U.S. patent application number 16/743982 was filed with the patent office on 2020-06-11 for process for leaching metal sulfides with reagents having thiocarbonyl functional groups.
The applicant listed for this patent is The University of British Columbia. Invention is credited to Edouard ASSELIN, David DIXON, Ahmad GHAHREMAN, Oscar OLVERA OLMEDO, Zihe REN.
Application Number | 20200181733 16/743982 |
Document ID | / |
Family ID | 57125622 |
Filed Date | 2020-06-11 |
View All Diagrams
United States Patent
Application |
20200181733 |
Kind Code |
A1 |
DIXON; David ; et
al. |
June 11, 2020 |
Process for Leaching Metal Sulfides with Reagents Having
Thiocarbonyl Functional Groups
Abstract
This application pertains to methods of recovering metals from
metal sulfides that involve contacting the metal sulfide with an
acidic sulfate solution containing ferric sulfate and a reagent
that has a thiocarbonyl functional group, wherein the concentration
of reagent in the acidic sulfate solution is sufficient to increase
the rate of metal ion extraction relative to an acidic sulfate
solution that does not contain the reagent, to produce a pregnant
solution containing the metal ions.
Inventors: |
DIXON; David; (Delta,
CA) ; OLVERA OLMEDO; Oscar; (Vancouver, CA) ;
ASSELIN; Edouard; (Richmond, CA) ; GHAHREMAN;
Ahmad; (Kingston, CA) ; REN; Zihe; (Vancouver,
CA) |
|
Applicant: |
Name |
City |
State |
Country |
Type |
The University of British Columbia |
Vancouver |
|
CA |
|
|
Family ID: |
57125622 |
Appl. No.: |
16/743982 |
Filed: |
January 15, 2020 |
Related U.S. Patent Documents
|
|
|
|
|
|
Application
Number |
Filing Date |
Patent Number |
|
|
15566684 |
Oct 13, 2017 |
|
|
|
PCT/CA2016/050444 |
Apr 15, 2016 |
|
|
|
16743982 |
|
|
|
|
62149015 |
Apr 17, 2015 |
|
|
|
Current U.S.
Class: |
1/1 |
Current CPC
Class: |
C22B 3/08 20130101; C01G
53/10 20130101; Y02P 10/234 20151101; C22B 23/043 20130101; Y02P
10/236 20151101; C22B 17/04 20130101; C22B 3/0004 20130101; C01G
49/14 20130101; C25C 1/12 20130101; C01G 11/00 20130101; C22B 3/42
20130101; C22B 15/0071 20130101; C01G 1/00 20130101; C01G 3/10
20130101; Y02P 10/20 20151101 |
International
Class: |
C22B 3/08 20060101
C22B003/08; C01G 1/00 20060101 C01G001/00; C01G 49/14 20060101
C01G049/14; C01G 11/00 20060101 C01G011/00; C01G 53/10 20060101
C01G053/10; C01G 3/10 20060101 C01G003/10; C25C 1/12 20060101
C25C001/12; C22B 3/20 20060101 C22B003/20; C22B 3/00 20060101
C22B003/00; C22B 15/00 20060101 C22B015/00; C22B 3/42 20060101
C22B003/42 |
Claims
1.-59. (canceled)
60. A method of recovering at least one base metal from at least
one base metal sulfide in a material, the method comprising:
contacting the material with an acidic sulfate solution comprising
a reagent having a thiocarbonyl functional group, wherein the
reagent is ethylene trithiocarbonate (ETC), to produce a pregnant
solution containing base metal ions; and recovering the at least
one base metal from the pregnant solution.
61. The method of claim 60, wherein the acidic sulfate solution
further comprises ferric sulfate.
62. The method of claim 60, wherein the concentration of the
reagent in the acidic sulfate solution is in the range of about 0.2
mM to about 30 mM.
63. The method of claim 60, wherein the concentration of the
reagent in the acidic sulfate solution is in the range of about 0.2
mM to about 20 mM.
64. The method of claim 60, wherein the concentration of the
reagent in the acidic sulfate solution is in the range of about 0.2
mM to about 10 mM.
65. The method of claim 60, wherein the concentration of the
reagent in the acidic sulfate solution is in the range of about 0.2
mM to about 5 mM.
66. The method of claim 60, wherein the concentration of the
reagent in the acidic sulfate solution is in the range of about 0.2
mM to about 4 mM.
67. The method of claim 60, wherein the concentration of the
reagent in the acidic sulfate solution is in the range of about 0.2
mM to about 3 mM.
68. The method of claim 60, wherein the concentration of the
reagent in the acidic sulfate solution is in the range of about 0.2
mM to about 2 mM.
69. The method of claim 60, wherein the concentration of the
reagent in the acidic sulfate solution is in the range of about 0.2
mM to about 1.5 mM.
70. The method of claim 60, wherein the concentration of the
reagent in the acidic sulfate solution is in the range of about 0.2
mM to about 1.0 mM.
71. The method of claim 60, wherein the concentration of the
reagent in the acidic sulfate solution is in the range of about 0.2
mM to about 0.5 mM.
72. The method of claim 60, wherein the at least one base metal
sulfide includes at least one copper sulfide.
73. The method of claim 72, wherein the at least one copper sulfide
includes chalcopyrite.
74. The method of claim 72, wherein the at least one copper sulfide
includes covellite.
75. The method of claim 72, wherein the at least one copper sulfide
includes bornite.
76. The method of claim 72, wherein the at least one copper sulfide
includes enargite.
77. The method of claim 72, wherein the at least one copper sulfide
includes a copper sulfide of the formula CuxSy wherein the x:y
ratio is between 1 and 2.
78. The method of claim 72, wherein the at least one copper sulfide
includes chalcocite.
79. The method of claim 72, wherein the at least one copper sulfide
includes djurleite.
80. The method of claim 72, wherein the at least one copper sulfide
includes digenite.
81. The method of claim 60, wherein the at least one base metal
sulfide includes a cadmium sulfide.
82. The method of claim 81, wherein the cadmium sulfide is
greenockite.
83. The method of claim 60, wherein the at least one base metal
sulfide includes at least one nickel sulfide.
84. The method of claim 83, wherein the at least one nickel sulfide
includes pentlandite.
85. The method of claim 83, wherein the at least one nickel sulfide
includes violarite.
86. The method of claim 60, wherein the material is an ore.
87. The method of claim 60, wherein the material is a concentrate
of the at least one base metal sulfide.
88. The method of claim 60, wherein the material comprises
agglomerated particles.
89. The method of claim 60, wherein ferric ions are used to oxidize
the metal sulfide.
90. The method of claim 89, wherein the ferric ions are generated
at least in part by bacteria.
91. The method of claim 60, wherein the method is a percolation
leach.
92. The method of claim 91, wherein the percolation leach is a heap
leach.
93. The method of claim 91, wherein the leach is a vat leach.
94. The method of claim 91, wherein the method is a tank leach.
95. The method of claim 91, wherein the method is a column
leach.
96. The method of claim 60, wherein recovering the at least one
base metal from the pregnant solution comprises solvent extraction
and electrowinning.
97. The method of claim 60, further comprising maintaining the
operating potential of the acidic sulfate solution above 500 mV vs
Ag/AgCl.
Description
[0001] This application claims priority to U.S. Patent Application
No. 62/149,015, filed Apr. 17, 2015, the contents of which are
incorporated herein by reference.
BACKGROUND
1. Field of the Disclosure
[0002] This disclosure pertains to methods for leaching metals from
metal sulfide-containing ores. More particularly it pertains to a
hydrometallurgical process for the extraction of metals from metal
sulfide-containing ores using reagents having a thiocarbonyl
functional group.
2. Description of Related Art
[0003] Aqueous processing of minerals presents several advantages
over pyrometallurgical approaches, particularly when dealing with
complex and/or low-grade ores. The main disadvantage of
hydrometallurgical processes, when applied to several metal sulfide
ores, is the low extraction rates that are observed. It is
desirable to develop a process where high metal extractions can be
achieved in time scales that are of industrial interest.
[0004] Chalcopyrite, for example, is a semiconductor, and therefore
corrodes electrochemically in oxidizing solutions. In ferric
sulfate media, the overall leaching reaction is as follows:
CuFeS.sub.2(s)+2Fe.sub.2(SO.sub.4).sub.3(a).fwdarw.CuSO.sub.4(a)+5FeSO.s-
ub.4(a)+2S.sup.0(s)
[0005] This reaction may be represented as a combination of anodic
and cathodic half-cell reactions:
CuFeS.sub.2.fwdarw.Cu.sup.2++Fe.sup.2++2S.sup.0+4e.sup.- Anodic
half-cell reaction:
4Fe.sup.3++4e.sup.-.fwdarw.4Fe.sup.2+ Cathodic half-cell
reaction:
[0006] A fundamental problem with chalcopyrite oxidation is that
chalcopyrite mineral surfaces become resistant to electrochemical
breakdown at solution potentials above a certain level (generally
considered to be about 550 to 600 mV vs Ag/AgCI). It is widely held
that this results from the formation of some sort of passivating
film on the mineral surface that most likely consists of an
altered, partially Fe-depleted form of chalcopyrite. It is
desirable to provide leaching processes in which such passivation
is reduced or avoided.
[0007] Some work has been done in extractive hydrometallurgy to
recover precious metals such as gold and silver from copper
concentrates or chalcopyrite residues after copper extraction.
Desch nes and Ghali (Hydrometallurgy 20:129-202) demonstrated the
potential application of thiourea in acidic sulfate leaching of
sulfide concentrates, such as those containing chalcopyrite, to
selectively recover gold and silver. Thiourea is an organosulfur
compound having a thiocarbonyl functional group. However, thiourea
did not appear to have an effect on the recovery of copper from
copper sulfides.
SUMMARY
[0008] This disclosure relates, at least in part, to the unexpected
discovery that several reagents comprising a thiocarbonyl function
group (e.g. thiourea) can be used to facilitate the leaching of
metal from several metal sulfides (e.g. copper from chalcopyrite)
with acidic sulfate leach solutions. When added in small amounts,
such reagents may increase the rate of metal leaching over that
observed in its absence.
[0009] This disclosure relates to a method of recovering at least
one metal from an ore containing at least one metal sulfide. The
method involves contacting the ore with an acidic sulfate solution
containing ferric sulfate and a reagent having a thiocarbonyl
functional group to extract metal ions from the at least one metal
sulfide, wherein the concentration of the reagent in the acidic
sulfate solution is sufficient to increase the rate of the metal
ion extraction relative to an acidic sulfate solution that does not
contain the reagent, to produce a pregnant solution containing the
metal ions. The method further involves recovering the at least one
metal from the pregnant solution. The at least one metal includes:
copper, wherein the at least one metal sulfide includes
chalcopyrite, covellite, bornite, enargite, a copper sulfide of the
formula Cu.sub.xS.sub.y wherein the x:y ratio is between 1 and 2,
or a combination thereof; cadmium, wherein the at least one metal
sulfide is greenockite; nickel, wherein the at least one metal
sulfide is pentlandite, violarite, or a combination thereof; or a
combination thereof. The concentration of the reagent may be in the
range of about 0.2 mM to about 30 mM.
[0010] This disclosure also relates to a method of recovering at
least one metal from an ore containing at least one metal sulfide.
The method involves contacting the metal sulfide with an acidic
sulfate solution containing a reagent having a thiocarbonyl
functional group, wherein the initial concentration of the reagent
in the acidic sulfate solution is in the range of about 0.2 mM to
about 30 mM or less, to produce a pregnant solution containing
metal ions. The method further involves recovering copper from the
pregnant solution. The at least one metal includes: copper, wherein
the at least one metal sulfide includes chalcopyrite, covellite,
bornite, enargite, a copper sulfide of the formula Cu.sub.xS.sub.y
wherein the x:y ratio is between 1 and 2, or a combination thereof;
cadmium, wherein the at least one metal sulfide is greenockite;
nickel, wherein the at least one metal sulfide is pentlandite,
violarite, or a combination thereof; or a combination thereof.
[0011] In the methods described above, the concentration of the
reagent in the acidic sulfate solution may be in the range of about
0.2 mM to about 20 mM, about 0.2 mM to about 10 mM, about 0.2 mM to
about 5 mM, about 0.2 mM to about 4 mM, about 0.2 mM to about 3 mM,
about 0.2 mM to about 2 mM, about 0.2 mM to about 1.5 mM, about 0.2
mM to about 1.0 mM, or about 0.2 mM to about 0.5 mM.
[0012] Where the metal is a copper sulfide of the formula CuxSy
wherein the x:y ratio is between 1 and 2, the copper sulfide may
includes chalcocite, djurleite, digenite, or a combination thereof.
In the methods described above, the reagent may be thiourea (Tu),
thioacetamide (TA), sodium-dimethyldithiocarbamate (SDDC), ethylene
trithiocarbonate (ETC), thiosemicarbazide (TSCA), or a combination
thereof.
[0013] This disclosure yet further relates to a method of
recovering a metal from an ore containing at least one metal
sulfide. The method involves contacting the ore with an acidic
sulfate solution comprising ferric sulfate and formamidine
disulfide (FDS) to produce a pregnant solution containing metal
ions. The method further involves recovering the metal from the
pregnant solution. The at least one metal includes: copper, wherein
the at least one metal sulfide includes chalcopyrite, covellite,
bornite, enargite, a copper sulfide of the formula Cu.sub.xS.sub.y
wherein the x:y ratio is between 1 and 2, or a combination thereof;
cadmium, wherein the at least one metal sulfide is greenockite;
nickel, wherein the at least one metal sulfide is pentlandite,
violarite, or a combination thereof; or a combination thereof. The
concentration of FDS in the acidic sulfate solution may be in the
range of about 0.1 mM to about 15 mM, about 0.1 mM to about 10 mM,
.about 0.2 mM to about 5 mM, about 0.1 mM to about 2.5 mM, about
0.1 mM to about 2 mM, about 0.1 mM to about 1.5 mM, about 0.1 mM to
about 1.0 mM, about 0.1 mM to about 0.5 mM, or about 0.1 mM to
about 0.25 mM. Where the metal is a copper sulfide of the formula
CuxSy wherein the x:y ratio is between 1 and 2, the copper sulfide
may includes chalcocite, djurleite, digenite, or a combination
thereof.
[0014] The concentration of FDS in the acidic sulfate solution may
be sufficient to provide sufficient thiourea to increase the rate
of the metal ion extraction relative to an acidic sulfate solution
that does not contain the reagent to produce the pregnant solution
containing the metal ions.
[0015] In the methods described above, wherein the ore may be
provided as coarse particles, which may be agglomerated particles.
Ferric ions may be used to oxidize the metal sulfide. In the
methods described above, the ferric ions may be generated at least
in part by bacteria.
[0016] The methods may involve a percolation leach. The percolation
leach may be a heap leach. The percolation leach may be a vat
leach. The leach may be a tank leach.
[0017] Recovering metal from the pregnant solution may include
solvent extraction and electrowinning.
[0018] Other aspects and features of the present invention will
become apparent to those ordinarily skilled in the art upon review
of the following description of specific embodiments of the
invention in conjunction with the accompanying figures.
BRIEF DESCRIPTION OF THE DRAWINGS
[0019] In drawings which illustrate embodiments of the
invention,
[0020] FIG. 1 is a plot showing the effect of thiourea
concentration on mixed potential and dissolution current density
(i.sub.dissol) of the CuFeS.sub.2 electrode;
[0021] FIG. 2 is a bar graph showing electrochemical dissolution
rates of a CuFeS.sub.2 electrode in sulfuric acid solution at pH 2
and 25.degree. C. with varying initial concentrations of thiourea,
formamidine disulfide (FDS), and Fe(III);
[0022] FIG. 3 is a schematic diagram for the leaching column used
in respect of the leaching experiments pertaining to FIGS. 4, 5,
and 6;
[0023] FIG. 4 is a graph showing the effect of thiourea
concentration on the leaching of copper from Ore A in column leach
experiments;
[0024] FIG. 5 is a graph showing the effect of thiourea
concentration on the leaching of copper from Ore B in column leach
experiments;
[0025] FIG. 6 is a graph showing the effect of thiourea
concentration on the leaching of copper from Ore C in column leach
experiments;
[0026] FIG. 7 is a graph showing the effect of thiourea
concentration on the leaching rate of copper from Ore C in column
leach experiments;
[0027] FIG. 8 is a graph showing the effect of thiourea
concentration on ORP over time;
[0028] FIG. 9 is a graph showing the effect of thiourea
concentration on copper dissolution for coarse Ore A in bottle roll
experiments;
[0029] FIG. 10 is a graph showing the effect of thiourea
concentration on copper dissolution for coarse Ore B in bottle roll
experiments;
[0030] FIG. 11 is a graph showing the effect of Tu addition on
various minerals that contain Cu(I). Diamonds pertain to bornite,
triangles refer to covellite, inverted triangles pertain to
chalcocite, and squares pertain to chalcopyrite. Open symbols refer
to control treatments without Tu, whereas solid symbols refer to
minerals treated solutions having an initial Tu concentration of 2
mM;
[0031] FIG. 12 is a graph showing the effect of Tu on cadium
extraction from greenockite;
[0032] FIG. 13 is a graph showing the effect of Tu on copper
extraction from enargite;
[0033] FIG. 14 is a graph showing the effect of Tu on nickel
extraction from violarite;
[0034] FIG. 15 is a graph showing the percentage of Cu ions
remaining in solution after various amounts of Tu addition;
[0035] FIG. 16 is a graph showing extraction of Cu from
chalcopyrite under various Tu dosages;
[0036] FIG. 17 is a graph showing the relationship between Tu
dosage and Cu extraction after 172 hours;
[0037] FIG. 18 is a graph showing leaching of copper from
chalcopyrite in stirred reactor tests using reagents comprising
thiocarbonyl functional groups. Circles pertain to Tu, triangles
pertain to TA, inverted triangles pertain to SDDC, diamonds pertain
to ETC, stars pertain to TSCA, and squares pertain to controls;
[0038] FIG. 19 is a graph showing leaching of copper from covellite
in stirred reactor tests using reagents comprising thiocarbonyl
functional groups. Circles pertain to Tu, triangles pertain to TA,
diamonds pertain to SDDC, and squares pertain to controls;
[0039] FIG. 20 is a graph showing leaching of copper from bornite
in stirred reactor tests using reagents comprising thiocarbonyl
functional groups. Triangles pertain to Tu, circles pertain to TA,
and squares pertain to controls;
[0040] FIG. 21 is a graph showing leaching of copper from enargite
in stirred reactor tests using reagents comprising thiocarbonyl
functional groups. Circles pertain to Tu, triangles pertain to TA,
inverted triangles pertain to ETC, and squares pertain to
controls;
[0041] FIG. 22 is a graph showing the leaching of copper from
chalcopyrite in stirred reactor tests using reagents comprising
thiocarbonyl functional groups, urea, and carbon disulfide. Circles
pertain to urea, triangles pertain to controls, inverted triangles
pertain to TA, diamonds pertain to Tu, stars pertain to ETC, and
squares pertain to carbon disulfide;
[0042] FIG. 23, panel a, is a graph comparing the leaching of
copper from chalcopyrite (circles) or bornite (triangles) using
leaching solutions with either an initial concentration of 2 mM Tu
(solid symbols) or an initial concentration of 1 m FDS (open
symbols);
[0043] FIG. 23, panel b, is a graph comparing the leaching of
copper from covellite (circles) or chalcocite (triangles) using
leaching solutions with either an initial concentration of 2 mM Tu
(solid symbols) or an initial concentration of 1 m FDS (open
symbols);
[0044] FIG. 24 is a graph monitoring bacterial activity and FDS
content with ORP and HPLC; and
[0045] FIG. 25 is a graph showing the bioleaching of CuFeS.sub.2
using only Fe.sup.3+ (day 0-50) and using Fe.sup.3++Tu (day 90-150)
in closed loop experiments.
DETAILED DESCRIPTION
[0046] This disclosure relates to methods of recovering metal from
a metal sulfide mineral, and relates in particular to the
unexpected discovery that various reagents having a thiocarbonyl
functional group, e.g. thiourea (also known as thiocarbamide), can
be used to facilitate the leaching of metal from a metal sulfide
mineral with acidic sulfate leach solutions. Such reagents can
increase the rate of metal sulfide leaching.
[0047] Such methods may be particularly useful in the recovery of
metal from low grade ores that do not contain the metal sulfide
mineral in high proportions. The method involves contacting the
copper sulfide mineral with an acidic sulfate solution containing
the reagent having a thiocarbonyl functional group.
Minerals
Chalcopyrite (CuFeS.sub.2)
[0048] The leaching of chalcopyrite is accomplished in acidic
ferric sulfate solution according to the following reaction
formula:
CuFeS.sub.2+4Fe.sup.3+.fwdarw.Cu.sup.2++5Fe.sup.2++2S.sup.0
Covellite (CuS)
[0049] Leaching of covellite in ferric sulfate solution proceeds
according to the following reaction formula:
CuS+2Fe.sup.3+.fwdarw.Cu.sup.2++2Fe.sup.2++S.sup.0
Chalcocite (Cu.sub.2S)
[0050] Leaching of chalcocite in ferric solution proceeds according
to the following formula:
Cu.sub.2S+2Fe.sup.3+.fwdarw.Cu.sup.2++2Fe.sup.2++CuS
[0051] The skilled person understands that that "chalcocite" ores
frequently contain a mixture of minerals with the formula
Cu.sub.xS.sub.y, where the x:y ratio is between 1 and 2. Additional
minerals within this formula include digenite and djurleite.
Bornite (Cu.sub.5FeS.sub.4)
[0052] Bornite is an important copper mineral that usually coexists
with chalcopyrite. The leaching process of bornite in ferric
solution is described in two stages:
Cu.sub.5FeS.sub.4+4Fe.sup.3+.fwdarw.Cu.sub.3FeS.sub.4+2Cu.sup.2++4Fe.sup-
.2+
Cu.sub.3FeS.sub.4+8Fe.sup.3+.fwdarw.3Cu.sup.2++9Fe.sup.2++4S.sup.0
Enargite (Cu.sub.3AsS.sub.4)
[0053] Unlike the other copper minerals mentioned above
(chalcopyrite, covellite, charcocite and bornite), the copper in
enargite is mainly Cu(II) instead of Cu(I). The difference in
copper's oxidation state will also influence its leaching kinetics
under catalyzed conditions. Previous study showed that the leaching
of enargite at atmospheric pressure is extremely slow. The
dissolution of enargite in ferric sulfate media can take various
paths. Two of them are described as follows:
Cu.sub.3AsS.sub.4+20H.sub.2O+35Fe.sup.3+.fwdarw.3Cu.sup.2++AsO.sub.4.sup-
.3-+4SO.sub.4.sup.2-+40H.sup.++35Fe.sup.2+
Cu.sub.3AsS.sub.4+4H.sub.2O+11Fe.sup.3+.fwdarw.3Cu.sup.2++AsO.sub.4.sup.-
3-+4S.sup.0+8H.sup.++11Fe.sup.2+
Greenockite (CdS)
[0054] Cadmium metal and compounds are mainly used for alloys,
coatings, batteries and plastic stabilizers. There are no mines
specifically for cadmium extraction. Cadmium sulfide is usually
associated with zinc sulfides and is recovered as a byproduct of
zinc leaching from roasted sulfide concentrates.
Violarite (FeNi.sub.2S.sub.4)
[0055] Violarite is a nickel (III) sulfide mineral that is usually
associated with primary pentlandite nickel sulfide ores.
Reagents
[0056] A person skilled in the art will also understand that
reagents having a thiocarbonyl functional group include, but are
not limited to thiourea (Tu), thioacetamide (TA),
sodium-dimethyldithiocarbamate (SDDC), ethylene trithiocarbonate
(ETC) and thiosemicarbazide (TSCA).
[0057] A non-exhaustive list of additional compounds having a
thiocarbonyl functional group is: isothiourea; N--N' substituted
thioureas; 2,5-dithiobiurea; dithiobiuret; Thiosemicarbazide purum,
Thiosemicarbazide; Methyl chlorothiolformate; Dithiooxamide;
Thioacetamide; 2-Methyl-3-thiosemicarbazide;
4-Methyl-3-thiosemicarbazide; Vinylene trithiocarbonate purum;
Vinylene trithiocarbonate; 2-Cyanothioacetamide; Ethylene
trithiocarbonate; Potassium ethyl xanthogenate;
Dimethylthiocarbamoyl chloride; dimethyldithiocarbamate;
S,S'-Dimethyl dithiocarbonate; Dimethyl trithiocarbonate;
N,N-Dimethylthioformamide; 4,4-Dimethyl-3-thiosemicarbazide;
4-Ethyl-3-thiosemicarbazide; O-Isopropylxanthic acid; Ethyl
thiooxamate; Ethyl dithioacetate; Pyrazine-2-thiocarboxamide;
Diethylthiocarbamoyl chloride; diethyldithiocarbamate;
Tetramethylthiuram monosulfide; Tetramethylthiuram disulfide;
Pentafluorophenyl chlorothionoformate; 4-Fluorophenyl
chlorothionoformate; O-Phenyl chlorothionoformate; O-Phenyl
chlorothionoformate; Phenyl chlorodithioformate;
3,4-Difluorothiobenzamide; 2-Bromothiobenzamide;
3-Bromothiobenzamide; 4-Bromothiobenzamide; 4-Chlorothiobenzamide;
4-Fluorothiobenzamide; Thiobenzoic acid; Thiobenzamide;
4-Phenylthiosemicarbazide; O-(p-Tolyl) chlorothionoformate;
4-Bromo-2-methylthiobenzamide; 3-Methoxythiobenzamide;
4-Methoxythiobenzamide; 4-Methylbenzenethioamide; Thioacetanilide;
Salicylaldehyde thiosemicarbazone; Indole-3-thiocarboxamide;
S-(Thiobenzoyl)thioglycolic acid; 3-(Acetoxy)thiobenzamide;
4-(Acetoxy)thiobenzamide; methyl
N'-[(e)-(4-chlorophenyl)methylidene]hydrazonothiocarbamate;
3-Ethoxythiobenzamide; 4-Ethylbenzene-1-thiocarboxamide; tert-Butyl
3-[(methylsulfonyl)oxy]-1-azetanecarboxylate; Diethyldithiocarbamic
acid; 2-(Phenylcarbonothioylthio)propanoic acid;
2-Hydroxybenzaldehyde N-ethylthiosemicarbazone;
(1R,4R)-1,7,7-Trimethylbicyclo[2.2.1]heptane-2-thione;
Tetraethylthiuram disulfide; Tetraethylthiuram disulfide;
4'-Hydroxybiphenyl-4-thiocarboxamide; 4-Biphenylthioamide;
Dithizone; 4'-Methylbiphenyl-4-thiocarboxamide;
tetraisopropylthiuram disulfide; Anthracene-9-thiocarboxamide;
Phenanthrene-9-thiocarboxamide; Sodium dibenzyldithiocarbamate; and
4,4'-Bis(dimethylamino)thiobenzophenone. Such agents are ready
available from, for example, Sigma Aldrich.
[0058] Each of Tu, TA, SDDC, ETC and TSCA feature a thiocarbonyl
functional group having a sulfur that 1) bears a partial negative
charge, 2) bears negative electrostatic potential surface, and 3)
has an empty .pi.*-antibonding orbital as its lowest unoccupied
molecular orbital (LUMO). Accordingly, the skilled person may
reasonably expect that other reagents, including those additional
reagents listed above, that share such criteria and are
sufficiently soluble in water may be useful in the performance of
the methods disclosed herein (provided that they do not complex
with the metal or iron oxidant to from precipitates). It will be
within the purview of the skilled person to identify potentially
useful reagents and test them to determine efficacy with any
particular ore, if any at all.
[0059] For example, Tu has a thiocarbonyl functional group with the
sulfur bearing a partial charge of -0.371, a negative electrostatic
potential around the Sulfur, and .pi.*-antibonding orbital as its
LUMO. Hence, thiourea satisfies all three criteria and has
demonstrated catalytic effect.
[0060] TA has a similar structure as Tu, but with a CH.sub.3 side
chain instead of NH.sub.2. It has a thiocarbonyl functional group
with the sulfur bearing a partial charge of -0.305, which is
slightly lower than that for Tu, a negative electrostatic potential
around the sulfur, and a .pi.*-antibonding orbital as its LUMO.
Accordingly, TA also satisfies all three criteria and has
demonstrated catalytic effect.
[0061] ETC differs from Tu and TA as it does not contain any
thioamide group. It has a thiocarbonyl functional group with the
two sulfur atoms .sigma.-bonded to carbon as the side chain. The
sulfur in the thiocarbonyl group bears a partial charge of -0.122,
which is much lower than Tu, a negative electrostatic potential
around the Sulfur, and .pi.*-antibonding orbital as its LUMO.
Accordingly, ETC also satisfies all three criteria and has
demonstrated catalytic effect.
[0062] In comparison, urea has a carbonyl functional group with a
C.dbd.O bond instead of C.dbd.S. The oxygen in the C.dbd.O bond
bears a partial charge of -0.634 and a negative electrostatic
potential around it, which is very similar to the sulfur atom in
Tu. However, its LUMO does not contain .pi.*-antibonding.
Accordingly, urea is not predicted to have a catalytic effect in
metal leaching, which is confirmed in respect of chalcopyrite by
the results of the stirred reactor experiment shown in FIG. 22.
[0063] Carbon disulfide (CS.sub.2) contains two thiocarbonyl
functional groups. Although the sulfur atoms of each functional
group contain a .pi.*-antibonding orbitals as their LUMO, they bear
a partial positive charge of +0.012. Therefore, CS.sub.2 is not
predicted to have catalytic effect, which is confirmed in respect
of chalcopyrite by the results of the stirred reactor experiment
shown in FIG. 23.
[0064] Of course, the reagent should also be water soluble. ETC,
for example, is only sparingly soluble in water, which may explain
why it appears less effective than Tu in leaching copper from
chalcopyrite.
[0065] Preferentially, the reagent will not form
complexes/precipitate with Fe.sup.2+/Fe.sup.3+ ions. TSCA, for
example, is able to form a red-color complex with Fe.sup.3+ in
solution, which may explain why it is less effective than Tu in
leaching copper from chalcopyrite.
[0066] The reagent also should not complex/precipitate with target
metal ions such as Cu.sup.+, Cu.sup.2+, Cd.sup.2+, or Ni.sup.2+.
Dithiooxamide forms an insoluble complex with copper ions and
therefore cannot be used for the leaching of copper sulfide
minerals, whereas TA complexes with Cd.sup.2+ ions to form an
insoluble complex and therefore cannot be used for leaching cadmium
sulfide minerals such as greenockite.
[0067] Again, the skilled person will appreciate that not all
compounds comprising a thiocarbonyl functional group will be useful
in increasing the rate of metal extraction from a metal sulfide.
Furthermore, the skilled person will appreciate that a reagent that
works to increase the rate of extraction of metal from one metal
sulfide may not be useful to increase the rate of extraction of a
metal from a different metal sulfide. Again, it will be within the
purview of the skilled person to identify potentially useful
reagents and test them to determine efficacy with any particular
ore, if any at all.
Formamidine Disulfide (FDS)
[0068] Formamidine disulfide (FDS) is generated by oxidation of Tu.
In the presence of an oxidant such as ferric sulfate, Tu will
oxidize partially to formamidine disulfide (FDS) according to the
following half-cell reaction:
2SC(NH.sub.2).sub.2.fwdarw.[(NH.sub.2).sub.2CS].sub.2.sup.2++2e.sup.-
[0069] FDS contains no thiocarbonyl functional group but a
sulfur-sulfur sigma bond instead. An equilibrium exists between FDS
and Tu in a ferric sulfate solution, such that a leach solution
prepared with FDS rather than Tu will provide the Tu necessary for
catalysis of the metal sulfide leach. That is, a molecule of FDS
will dissociate into two molecules of Tu upon dissolution in the
ferric sulfate leach solution. Accordingly, a leaching solution
employing Tu as the reagent having the thiocarbonyl functional
group may be effectively be prepared using either Tu or FDS.
[0070] The skilled person will understand that, due to this
equilibrium, the concentration of Tu (and FDS) may fluctuate over
time. Accordingly, "concentration" as used herein to refer to the
concentration of Tu in the leach solution relates to the amount of
Tu present in the solution as if all FDS in the solution was
dissociated into Tu (i.e ignoring interconversion between the two
forms). Similarly, "concentration" as used herein to refer to the
concentration of FDS in the leach solution relates to the amount of
FDS present in the solution as if all Tu in the solution was
converted into FDS (i.e ignoring interconversion between the two
forms).
[0071] "Initial concentration" is used herein to refer to the
initial concentration of the reagent at the time the leach solution
is applied to the ore sample. However, the skilled person will
understand that the concentration of the reagent may diminish over
time (e.g. through precipitation or decay) as the solution
percolates through the column or the heap. Accordingly, the skilled
person will appreciate that the processes disclosed herein should
work to increase the rate of metal extraction from the metal
sulfide provided that the concentration of the reagent is within a
suitable range during some portion of the percolation through the
ore.
[0072] In the presence of FDS and ferric sulphate (or another
suitable oxidant), the anodic dissolution of a copper sulfide
mineral such as chalcopyrite may proceed according to the following
two reactions, with oxidation of the chalcopyrite by either FDS or
ferric, respectively:
CuFeS.sub.2(s)+2[(NH.sub.2).sub.2CS].sub.2SO.sub.4(aq).fwdarw.CuSO.sub.4-
(aq)+FeSO.sub.4(aq)+2S.sup.0(s)+4SC(NH.sub.2).sub.2(aq)
CuFeS.sub.2(s)+2Fe.sub.2(SO.sub.4).sub.3(a).fwdarw.CuSO.sub.4(a)+5FeSO.s-
ub.4(a)+2S.sup.0(s)
[0073] After chalcopyrite is oxidized, and the copper is leached
from the concentrate, it is desirable to recover the copper from
the pregnant leach solution.
[0074] The methods disclosed herein involve two basic steps,
namely, leaching and metal recovery (e.g. by SX-EW). The leaching
process may be carried out as a percolation leach (such as a heap
leach), a vat leach, or a tank leach as is known in the field.
[0075] For the purposes of this disclosure, the words "containing"
and "comprising" are used in a non-limiting sense to mean that
items following the word are included, but items not specifically
mentioned are not excluded. A reference to an element by the
indefinite article "a" does not exclude the possibility that more
than one of the elements is present, unless the context clearly
requires that there be one and only one of these elements.
[0076] A "percolation leach", as used herein, refers to the
selective removal of a mineral by causing a suitable solvent to
seep into and through a mass or pile of material containing the
desired soluble mineral, e.g. a column leach or a heap leach.
[0077] A "column leach", as used herein, refers to leaching through
the use of a long narrow column in which ore sample and solution
are in contact for measuring the effects of typical variables
encountered in actual heap leaching.
[0078] A "heap leach", as used herein, is a process through which
metals are extracted from the ore in which they are found, i.e.
without beneficiation. A heap leach is often chosen for its
efficiency and cost-effectiveness. After being removed from the
ground, ore is typically sent through a crusher to break the ore
down into smaller particles (although heap ores can be
"run-of-mine" in which the ore is leached in an "as-blasted" state
with no further crushing). Heap ores may be the product of primary,
secondary, or tertiary crushing. Traditionally, the crushed
particles are then "heaped", or "stacked" into a large pile.
[0079] A persistent cause of failure of heap leach operations is
the presence of excess fines in the materials placed on the pad.
Excess fines results in a low permeability material and thus the
seepage rate of the lixiviant is too slow, or ore-solution contact
is insufficient, for economic pad operations. Accordingly, the
efficiency of a heap leach may be increased by agglomeration after
crushing. "Agglomeration", as used herein, refers to a technique
that binds together material fines or particles to create a larger
product. Agglomeration may be achieved by different methods known
in the art. Typically, heap leach agglomeration is performed in a
drum agglomerator with sulfuric acid and no binder, or on conveyor
belts with acid sprayed onto the ore at drop points.
[0080] The heap is irrigated with a solution that is dependent upon
the type of ore being extracted. Acid for the leach will preferably
be generated by bacteria using processes known in the art.
Alternatively, additional acid could be added as necessary.
[0081] The irrigated solution is allowed to percolate through the
ore, and drain to the bottom of the heap. The ore pile sits over an
impermeable layer, such as plastic sheet, which collects the
pregnant leach solution as it drains through and directs it to a
collection pond. Once the solution is collected, it is pumped to a
recovery plant to extract the copper by solvent extraction and
electrowinning (SX-EW).
[0082] Applying the methods disclosed herein to a heap leach, ore
containing an appropriate sulfide mineral is leached selectively in
the presence of the acid sulfate and the reagent having a
thiocarbonyl functional group. The concentration of the reagent
having a thiocarbonyl functional group in the leach solution may be
about 30 mM or perhaps even higher. The skilled person will
understand that it is only necessary that the reagent concentration
be within a range sufficient to increase the leach rate of the
metal sulfide.
[0083] Moreover, while the results presented herein indicate that
reagent concentrations of about 30 mM or less are sufficiently low
to facilitate the leaching of metal from a particular metal
sulfide, 30 mM concentrations may not be economically feasible at
the present time. Accordingly, it may be preferable to use lower
concentrations of reagent that are feasible from economic and
operational points of view, e.g. about 20 mM or less, about 10 mM
or less, about 5 mM or less, about 4 mM or less, about 3 mM or
less, about 2 mM or less, about 1.5 mM or less, about 1 mM or less,
about 0.9 mM or less, about 0.8 mM or less, about 0.7 mM or less,
about 0.6 mM or less, about 0.5 mM or less, about 0.4 mM or less,
0.3 mM or less, or about 0.2 mM.
[0084] Accordingly, the concentration of the reagent in the acidic
sulfate solution may in the range of about 0.2 mM to about 0.3 mM,
about 0.2 mM to about 0.4 mM, about 0.2 mM to about 0.5 mM, about
0.2 mM to about 0.6 mM, about 0.2 mM to about 0.7 mM, about 0.2 mM
to about 0.8 mM, about 0.2 mM to about 0.9 mM, about 0.2 mM to
about 1.0 mM, about 0.2 to about 1.5 mM, about 0.2 to about 2.0 mM,
about 0.2 to about 2.5 mM, about 0.2 to about 3 mM, about 0.2 to
about 4 mM, about 0.2 to about 5 mM, about 0.2 to about 10 mM,
about 0.2 to about 20 mM, or about 0.2 to about 30 mM.
[0085] The leaching process may be run at temperatures between
0.degree. C. (i.e. the freezing point of water) and 80.degree. C.
However, the process would typically be carried out at ambient
temperature and atmospheric pressure.
[0086] Following the leaching process, copper can be extracted from
the leach solution. After a solid-liquid separation, i.e. drainage
of the pregnant leach solution containing the copper from the heap,
the pregnant solution is preferably subjected to conventional
solvent extraction and electrowinning to produce pure copper
cathodes according to the following overall reaction:
SX-EW:
CuSO.sub.4(a)+H.sub.2O(l).fwdarw.Cu(s)+H.sub.2SO.sub.4(a)+1/2O.su-
b.2(g)
[0087] Reagents having a thiocarbonyl functional group in the
pregnant leach solution should not present any problem in the
electrowinning operation and, as a matter of fact, may even be
useful as a leveling agent. Raffinate containing thiourea may then
be recirculated to the heap for further leaching. The recirculated
leach solution may also be supplemented with thiourea to arrive at
the desired initial thiourea concentration for the leach.
EXAMPLES
[0088] To facilitate the extraction of metal ions from the minerals
listed above, reagents having a thiocarbonyl functional group were
added to acidic ferric sulfate solutions as catalysts. In the
experiments disclosed herein, it was found that the reagents that
contain thiocarbonyl functional groups have positive catalytic
effect on the extraction of the minerals. Among all the reagents,
Tu consistently provided the highest catalytic performance.
Accordingly, Tu was the most heavily studied reagent of those
identified. However, the results of experiments with other reagents
having thiocarbonyl functional groups are provided to compare their
catalytic effects. FDS which does not contain a thiocarbonyl
functional group but has comparable catalytic effect as Tu was
studied as a special case due to its equilibrium with Tu.
[0089] Leaching reactions were carried out at atmospheric pressure
on a variety of ore compositions, reagent concentrations, ferric
concentrations, and under various other conditions, as described
below.
Example 1 Extraction of Copper from Chalcopyrite Using Thiourea
Example 1.1
[0090] The effect of thiourea on the electrochemical behavior of a
chalcopyrite electrode was studied in a conventional 3-electrode
glass-jacketed cell. A CuFeS.sub.2 electrode was using as working
electrode, a saturated calomel electrode (SCE) was used as
reference, and a graphite bar was used as counter-electrode. The
CuFeS.sub.2 electrode was polished using 600 and 1200 grit carbide
paper. All experiments were conducted at 25.degree. C. using a
controlled temperature water bath. The electrolyte composition was
500 mM H.sub.2SO.sub.4, 20 mM Fe.sub.2SO.sub.4 and 0-100 mM
thiourea. Before starting any measurement, solutions were bubbled
with N.sub.2 for 30 minutes to reduce the concentration of
dissolved O.sub.2. Open circuit potential (OCP) was recorded until
changes of no more than 0.1 mV/min were observed. After a steady
OCP value was observed, electrochemical impedance spectroscopy
(EIS) was conducted at OCP using a 5 mV a.c. sinusoidal
perturbation from 10 kHz to 10 mHz. Linear polarization resistance
(LPR) tests were also conducted using a scan rate of 0.05 mV/s at
.+-.15 mV from OCP.
[0091] Linear potential scans were conducted at electrode
potentials .+-.15 mV from the OCP measured at each thiourea
concentration. All scans showed a linear behavior within the
electrode potential range analyzed. An increase in the slope of the
experimental plots was observed with increasing thiourea
concentration. The slope of these curves was used to estimate the
value of the polarization resistance (R.sub.ct) at each
concentration. These values were then used to estimate the values
of the dissolution current density using equation 1:
i dissol .apprxeq. RT nFR ct Eq . ( 1 ) ##EQU00001##
[0092] FIG. 1 shows the effect of thiourea on the dissolution
current density and mixed potential of the CuFeS.sub.2 electrode,
and indicates that a maximum dissolution current density was
achieved when thiourea concentration is 30 mM. Increasing thiourea
concentration to 100 mM resulted in a decrease in the current
density and mixed potential of the CuFeS.sub.2 electrode. Moreover,
after immersing the CuFeS.sub.2 electrode in the 100 mM thiourea
solution, a copper-like film was observed on the surface of the
electrode, which film could only be removed by polishing the
electrode with carbide paper.
Example 1.2
[0093] FIG. 2 is a bar graph showing the effect of initial thiourea
or FDS concentration on the electrochemical dissolution of a
chalcopyrite electrode in sulfuric acid solution at pH 2 and
25.degree. C. A concentration of 10 mM thiourea in the leach
solution resulted in a six fold increase in dissolution rate
compared to no thiourea, and a concentration of 5 mM FDS resulted
in a six fold increase relative to 10 mM thiourea. A concentration
of 10 mM thiourea in leach solution also containing 40 mM Fe(III)
resulted in a thirty fold increase in dissolution rate compared to
40 mM Fe(III) alone.
Example 1.3
[0094] A column leach of different acid-cured copper ores was
conducted with thiourea added to the leach solution. A schematic
description of the column setup is shown in FIG. 3. The column
diameter was 8.84 cm, the column height was 21.6 cm, and the column
stack height was 15.9 cm. The irrigation rate was 0.77 mL/min or 8
L/m.sup.2/h. The pregnant leach solution emitted from these columns
was sampled for copper every 2 or 3 days using Atomic Absorption
Spectroscopy (AAS).
[0095] The specific mineralogical composition of these ores are
provided in Table 1. The Cu contents of Ore A, Ore B, and Ore C
were 0.52%, 1.03%, and 1.22% w/w, respectively. Prior to leaching,
ore was "acid cured" to neutralize the acid-consuming material
present in the ore. That is, the ore was mixed with a concentrated
sulfuric acid solution composed of 80% concentrated sulfuric acid
and 20% de-ionized water and allowed to sit for 72 hours. For one
treatment using Ore C, thiourea was added to the sulfuric acid
curing solutions.
[0096] The initial composition of the leaching solutions included
2.2 g/L Fe (i.e. 40 mM, provided as ferric sulfate) and pH 2 for
the control experiment, with or without 0.76 g/L thiourea (i.e. 10
mM). The initial load of mineral in each column was 1.6 to 1.8 kg
of ore. The superficial velocity of solution through the ore column
was 7.4 L m.sup.-2 h.sup.-1. The pH was adjusted using diluted
sulfuric acid. These two columns were maintained in an open-loop or
open cycle configuration (i.e. no solution recycle) for the entire
leaching period.
[0097] The results of leaching tests on the Ore A, Ore B and Ore C
are shown in FIGS. 4, 5, and 6, respectively. The presence of
thiourea in the lixiviant clearly has a positive effect on the
leaching of copper from the chalcopyrite. On average, the leaching
rate in the presence of thiourea was increased by a factor of 1.5
to 2.4 compared to the control tests in which the leach solutions
did not contain thiourea. As of the last time points depicted in
FIGS. 4 to 6, copper extractions for columns containing Ore A, Ore
B, and Ore C leached with a solution containing sulfuric acid and
ferric sulfate alone, without added thiourea, were 21.2% (after 198
days), 12.4% (after 50 days), and 40.6% (after 322 days),
respectively. With 10 mM of added thiourea, these extractions were
37.9%, 32.0%, and 72.3%, respectively.
[0098] Referring to FIG. 6, 2 mM Tu was added to the leach solution
originally containing no Tu from day 322 onward, after which the
leach rate increased sharply. From day 332 to day 448, the copper
leached from this column increased from 40% to 58%, and rapid
leaching was maintained throughout that period.
[0099] The averages for the last 7 days reported in FIG. 7 indicate
that the leaching rate for acid-cured Ore C leached in the presence
of 10 mM thiourea is 3.3 higher than for acid-cured Ore C leached
in the absence of thiourea, and 4.0 times higher than acid-cured
and thiourea-cured Ore C leached in the absence of thioruea.
[0100] FIG. 8 shows the effect of Tu on solution potential. All
potentials are reported against a Ag/AgCl (saturated) reference
electrode. The solution potential of the leach solutions containing
thiourea was generally between 75 and 100 mV lower than the
solution potential of leach solution that did not include thiourea.
Lower solution potentials are consistent with thiourea working to
prevent the passivation of chalcopyrite.
Example 1.4 Bottle Roll Leaching
[0101] "Bottle roll" leaching experiments in the presence of
various concentrations of thiourea were conducted for coarse Ore A
and Ore B. The tests were conducted using coarsely crushed (100%
passing 1/2 inch) ore.
[0102] Prior to leaching, the ore was cured using a procedure
similar to what was performed on the ore used in the column
leaching experiments. The ore was mixed with a concentrated
sulfuric acid solution composed of 80% concentrated sulfuric acid
and 20% de-ionized water and allowed to settle for 72 hours to
neutralize the acid-consuming material present in the ore. For
several experiments, different concentrations of thiourea were
added to the ore using the sulfuric acid curing solutions.
[0103] The bottles used for the experiments were 20 cm long and
12.5 cm in diameter. Each bottle was loaded with 180 g of cured ore
and 420 g of leaching solution, filling up to around one third of
the bottle's volume.
[0104] The leaching solution from each bottle was sampled at 2, 4,
6 and 8 hours, and then every 24 hours thereafter. Samples were
analyzed using atomic absorption spectroscopy (AAS) for their
copper content.
[0105] The conditions for the bottle roll experiments are listed in
Table 2. Experiments #1 to #6 were conducted using only the
original addition of thiourea into the bottles. For experiments #7
to #11, thiourea was added every 24 hours to re-establish the
thiourea concentration.
[0106] A positive effect of thiourea on copper leaching was
observed. For the coarse ore experiments, a plateau was not
observed until after 80 to 120 hours. Thiourea was added
periodically to the coarse ore experiments, yielding positive
results on copper dissolution.
[0107] The effect of different concentrations of thiourea in the
leach solution on the leaching of coarse ore (experiments #1 to #11
as described in Table 2) is shown in FIGS. 9 and 10.
[0108] For ore B, thiourea was periodically added every 24 hours to
re-establish the thioruea concentration in the system and thus
better emulate the conditions in the column leach experiments. As
may be observed from FIG. 9, 8 mM and 10 mM thiourea yielded higher
copper dissolution results than the other thiourea concentrations
tested for ore A. A plateau in dissolution is not observed until
after approximately 120 hours, which varied with thiourea
concentration as shown in FIG. 9.
TABLE-US-00001 TABLE 1 Mineral Ideal Formula Ore A Ore B Ore C
Actinolite Ca.sub.2(Mg, Fe.sup.2+).sub.5Si.sub.8O.sub.22(OH).sub.2
-- 1.8 -- Biotite K(Mg,
Fe.sup.2+).sub.3AlSi.sub.3O.sub.10(OH).sub.2 -- 4.2 -- Calcite
CaCO.sub.3 -- 19.3 -- Chalcopyrite CuFeS.sub.2 1.4 3.5 2.6
Clinochlore (Mg, Fe.sup.2+).sub.5Al(Si.sub.3Al)O.sub.10(OH).sub.8
-- 15.0 -- Diopside CaMgSi.sub.2O.sub.6 -- 3.5 -- Galena PbS -- --
0.1 Gypsum CaSO.sub.42H.sub.2O -- 1.2 -- Hematite
.alpha.-Fe.sub.2O.sub.3 -- 0.2 -- K-feldspar KAlSi.sub.3O.sub.8
17.9 10.8 -- Kaolinite Al.sub.2Si.sub.2O.sub.5(OH).sub.4 2.3 -- 2.3
Magnetite Fe.sub.3O.sub.4 -- 0.8 -- Molybdenite MoS.sub.2 <0.1
-- -- Muscovite KAl.sub.2A1Si.sub.3O.sub.10(OH).sub.2 21.9 6.0 41.6
Plagioclase NaAlSi.sub.3O.sub.8--CaAlSi.sub.2O.sub.8 13.6 25.4 --
Pyrite FeS.sub.2 2.3 -- 8.0 Quartz SiO.sub.2 40.0 8.3 44.4 Rutile
TiO.sub.2 0.5 -- 0.9 Siderite Fe.sup.2+CO.sub.3 -- 0.1 -- Total 100
100 100
[0109] As may be observed from FIG. 9, 5 mM thiourea yielded higher
copper dissolution results than the other thiourea concentrations
tested for ore B. As with ore A, a plateau in dissolution is not
observed until after approximately 80 to 120 hours, which varied
with thiourea concentration as shown in FIG. 9. Periodic addition
of thiourea resulted in increased copper dissolutions and produced
a delay in the dissolution plateau.
[0110] Interestingly, solutions containing 100 mM thiourea did not
appear to be much more effective on copper extraction than those
containing no thiourea, and even worse at some time points. This is
consistent with the results of Desch nes and Ghali, which reported
that solutions containing .about.200 mM thiourea (i.e. 15 g/L) did
not improve copper extraction from chalcopyrite. Thiourea is less
stable at high concentrations and decomposes. Accordingly, it is
possible that, when initial thiourea concentrations are somewhat
higher than 30 mM, sufficient elemental sulfur may be produced by
decomposition of thiourea to form a film on the chalcopyrite
mineral and thereby assist in its passivation. It is also possible
that, at high Tu dosages, some copper precipitates from solution
(e.g. see FIG. 15) to account for some of the low extraction
results.
Example 2 Extraction from Chalcopyrite, Covellite, Chalcocite,
Bornite, Enargite, Pentlandite, Violarite, and Greenockite Using
Thiourea
[0111] The catalytic effect of Tu was further demonstrated in
stirred reactor tests. All reactors contained 1.9 L of ferric
sulfate solution at pH 1.8 and total iron concentration of 40 mM. 1
g of mineral samples was used in each reactor test. These
experimental conditions were designed to maintain an unlimited
supply of oxidant.
[0112] In order to demonstrate the catalytic effect on
chalcopyrite, 100% pure synthetic chalcopyrite was used instead of
chalcopyrite concentrate which contains various impurities. The
chalcopyrite was synthesized via a hydrothermal approach. CuCl,
FeCl.sub.3 and thiourea were first mixed with a molar ratio of
1:1:2 and dissolved in 150 mL DI water. The solution was
transferred to a Teflon-lined reaction vessel and heated up to
240.degree. C. for 24 hours. At the end of the reaction, the
precipitated powder was washed with acidic water (pH=1) and dried
in air at room temperature. XRD analysis in showed that the
synthetic chalcopyrite was free of any impurities compared with
chalcopyrite mineral concentrate. This synthetic chalcopyrite was
used in all the tests carried out in stirred reactors as disclosed
herein.
[0113] The covellite mineral used in the experiment disclosed
herein was also synthesized via a hydrothermal approach. CuCl and
Tu were mixed with a molar ratio of 1:1 and dissolved in 150 mL DI
water. The solution was transferred to a Teflon-lined reaction
vessel and heated up to 220.degree. C. for 24 hours. The
synthesized CuS was acid-washed and dried in air. XRD analysis
showed that it had 100% purity with no interference of other
species.
[0114] The chalcocite mineral sample used in the experiments
disclosed herein was 100% pure natural mineral.
[0115] The bornite mineral used in the experiments disclosed herein
was obtained from Butte, Mont. with copper content of 58.9% based
on ICP-AES. XRD analysis showed that the mineral contains 76.8%
bornite, 8.1% chalcopyrite, 6.3% pyrite, 5.8% tennatite and 3.0%
enargite. The copper content calculated from XRD was 55.6%, which
is relatively consistent with the chemical assay.
[0116] The enargite mineral used in the experiments disclosed
herein was in the form of an enargite concentrate, which contained
approximately 70% enargite (34% copper) according to XRD
analysis.
[0117] The Greenockite mineral used in this experiment was
synthesized via a hydrothermal approach. CdCl.sub.2 and thiourea
were mixed with a molar ratio of 1:1 and dissolved in 100 mL DI
water. The solution was transferred to a Teflon-lined reaction
vessel and heated up to 150.degree. C. for 24 hours. The
synthesized CdS was acid-washed and dried in air. XRD analysis
showed that it has 100% purity with no interference of other
species.
TABLE-US-00002 TABLE 2 List of bottle roll leaching experiments
involving Ore A and Ore B. Experiment Brief description of
experimental conditions #1 Coarse ore A, 0 mM Tu in solution, 40 mM
ferric in solution, acid curing, no Tu replenishment #2 Coarse ore
A, 2 mM Tu in solution, 40 mM ferric in solution, acid curing, no
Tu replenishment #3 Coarse ore A, 4 mM Tu in solution, 40 mM ferric
in solution, acid curing, no Tu replenishment #4 Coarse ore A, 6 mM
Tu in solution, 40 mM ferric in solution, acid curing, no Tu
replenishment #5 Coarse ore A, 8 mM Tu in solution, 40 mM ferric in
solution, acid curing, no Tu replenishment #6 Coarse ore A, 10 mM
Tu in solution, 40 mM ferric in solution, acid curing, no Tu
replenishment #7 Coarse ore B, 0 mM Tu in solution, 40 mM ferric in
solution, acid curing #8 Coarse ore B, 1 mM Tu in solution, 40 mM
ferric in solution, acid curing, periodic addition of Tu to
replenish 1 mM concentration in solution #9 Coarse ore B, 5 mM Tu
in solution, 40 mM ferric in solution, acid curing, periodic
addition of Tu to replenish 5 mM concentration in solution #10
Coarse ore B, 10 mM Tu in solution, 40 mM ferric in solution, acid
curing, periodic addition of Tu to replenish 10 mM concentration in
solution #11 Coarse ore B, 100 mM Tu in solution, 40 mM ferric in
solution, acid curing, periodic addition of Tu to replenish 100 mM
concentration in solution
[0118] The violarite used in the experiments disclosed herein was
natural violarite mineral that contains 15.8% Ni according to
ICP-AES. XRD analysis showed that the mineral had approximately 42%
violarite and 13.1% NiSO.sub.4.6H.sub.2O.
[0119] The sulfur on thiocarbonyl groups contains a lone electron
pair and a filled .pi.-orbital which can be used for donor-acceptor
type bonding with a transition metal, together with a
.pi.*-antibonding orbital that could potentially accept the
back-donation of electrons from the filled d-orbitals on the
transition metal. Accordingly, without wanting to be bound by
theory, it is suspected that the interaction between the surface
ion and the thiocarbonyl functional group, especially back donation
from metal to ligand, is responsible for the catalytic effect.
Moreover, it is suspected that the catalytic effect should be more
pronounced for the transition metals with higher d-electron
numbers, with the catalytic effect being most pronounced for
minerals with d.sup.10 electronic configuration
[0120] FIG. 11 shows that Tu catalyzes the leaching of common
copper sulfide minerals, including chalcopyrite, covellite,
chalcocite, and bornite, which all contain Cu(I). After 96 hours of
leaching, chalcopyrite extraction reaches 64.1% with 2 mM of Tu
compared to 21.1% without Tu; covellite extraction reaches 74.4%
with 2 mM of Tu compared to 7.2% without Tu; chalcocite extraction
reaches 85.6% with 2 mM of Tu compared to 65.1% without Tu; bornite
extraction reaches 91.4% with 2 mM of Tu compared to 56.7% without
Tu.
[0121] Like Cu(I), Cd(II) also contains the d.sup.10 electronic
configuration. FIG. 12 shows that leaching of CdS mineral is
significantly enhanced with the addition of Tu. With Tu, the
extraction of cadmium reaches 100% at 48 hours whereas extraction
in the noncatalyzed reaction plateaued at 47% after 96 hours.
[0122] The copper ion in the enargite mineral has fewer d-electrons
than other primary and secondary sulfides, and thus it may be
expected that the catalytic effect should be slower than that for
Cu(I) minerals. Nevertheless, the results shown in FIG. 13 clearly
demonstrate that leaching with a leach solution comprising an
initial concentration of 2 mM Tu increases the leach rate of copper
from enargite compared to a control without Tu, which did not show
any significant extraction after 96 hours of leaching.
[0123] Minerals that contain transition metal ions with d.sup.7
electronic configuration, such as Ni(III), may also undergo
catalyzed leaching with the addition of Tu. Similar to Cu(II), as
Ni(III) is the highest stable oxidation state with 7 d-electrons,
the catalytic effect is not expected to be as dramatic as for
d.sup.10 minerals. Referring to FIG. 14, leaching with a leach
solution comprising an initial concentration of 2 mM Tu increases
the leach rate of nickel from violarite compared to a control
without Tu.
[0124] Results of leaching experiments referred to in Example 2 are
summarized in Table 3, in which the extraction percentages under
non-catalyzed and catalyzed conditions (with an initial Tu
concentration of 2 mM) are compared.
TABLE-US-00003 TABLE 3 Comparisons of reactor leaching for various
minerals under uncatalyzed and 2 mM Tu catalyzed conditions 96-Hour
96-Hour Extraction Extraction Mineral (No thiourea) (2 mM thiourea)
Chalcopyrite, CuFeS.sub.2 21.1% 64.1% Covellite, CuS 6.8% 74.4%
Chalcocite, Cu.sub.2S 65.1% 85.5% Bornite, Cu.sub.5FeS.sub.4 56.7%
91.4% Greenokite, CdS 46.5% 100.0% Enargite, Cu.sub.3AsS.sub.4 2.1%
10.0% Violarite, FeNi.sub.2S.sub.4 13.0% 22.2%
Example 3 Reagent Dosage
[0125] Optimum dosage of reagent may increase the efficiency of
leaching. First, at certain concentrations, the reagent may form an
insoluble complex with the metal ion of interest and precipitate.
For example, Tu can form an insoluble complex with Cu(I) ions at a
3:1 molar ratio. A precipitation test was performed to examine the
concentration range at which Cu-Tu complex precipitation may occur.
20 mL of Cu solution was divided into several identical portions
followed by the addition of various Tu dosage (i.e. 0 to 60 mM).
The solution was stirred for 24 hours, and the Cu remaining in the
solution phase was analyzed by AAS. The results are shown in FIG.
15, plotted as the percentage of Cu remaining.
[0126] Second, heap leaching of metal sulfides is based on a
bioleaching mechanism, an excessive amount of reagent may be
detrimental to bioleaching microbes. For example, bacteria commonly
used for bioleaching, such as Acidithiobacillus ferrooxidans and
Acidithiobacillus thiooxidarns, have very slow growth in a solution
containing 10 mM Tu, and cannot survive at 100 mM Tu.
[0127] Third, with respect to Tu specifically, ferric reacts with
Tu and converts it to FDS (see Hydrometallurgy 28, 381-397 (1992)).
Although the reaction is reversible under certain conditions, a
high concentration of FDS tends to decompose irreversibly into
cyanamide and elemental sulfur (see J Chromatogr 368, 444-449).
2Tu+2Fe.sup.3+FDS+2Fe.sup.2++2H.sup.+
FDS.fwdarw.Tu+cyanimide+S
[0128] Therefore, over-addition of Tu in the lixiviant may cause
the loss of Fe.sup.3+ and Tu due to oxidation and decomposition.
The irreversible decomposition of FDS has been observed when adding
4 mM of Tu into a 40 mM ferric sulfate solution at pH 1.8.
[0129] To further investigate the effect of Tu dosage on copper
extraction, stirred reactor tests were performed using 1 g of
synthetic chalcopyrite in 1.9 L of 40 mM ferric sulfate solution at
pH 1.8 with various initial Tu concentrations. The treatments were
run for 172 hours to approach maximum extraction. The results are
presented in FIG. 16, and shows that, for 1 g of chalcopyrite,
higher Tu dosage results in faster leaching kinetics among the Tu
concentrations tested.
[0130] For Tu dosages of 5 mM and under, the initial 40 mM ferric
sulfate solution can be considered as a sufficient supply of
oxidant. However, for higher dosages such as 10 mM and 20 mM of Tu,
extra ferric (in 1:1 ratio with Tu) had to be added to the solution
to allow the oxidation of Tu to FDS. For 10 mM Tu, an extra 10 mM
of Fe.sup.3+ was added at time zero. For 20 mM Tu, an extra 20 mM
of Fe.sup.3+ was added at 72 hours, which led to the continuation
of extraction as shown in FIG. 16.
[0131] The Tu dosage vs. Cu extraction at 172 hours is plotted in
FIG. 17. An initial Tu dosage up to 5 mM appears to have the most
pronounced effect on the dissolution of Cu.
[0132] As indicated above, in previous shakeflask tests with acidic
solutions (pH 1.8) containing various concentrations of Fe.sup.3+
and Cu.sup.2+ ions, slight precipitation occurred upon the addition
of 4 mM of Tu due to the decomposition of FDS. Accordingly,
concentrations of Tu concentration below 4 mM may avoid such
precipitation. A series of shakeflask tests were performed on
solutions containing initial concentrations of 2 mM Tu and various
concentrations in a matrix containing Fe.sup.3+ (0-100 mM) and
Cu.sup.2+ (0-50 mM) in order to identify concentration ranges of
[Fe.sup.3+] and [Cu.sup.2+] that do not result in Cu complex
precipitation. The results showed that no precipitation and no loss
of Cu from the solution phase resulted using 2 mM of Tu in this
wide range of Fe and Cu matrix concentrations.
Example 4 Alternative Reagents
[0133] The catalytic effect of several other reagents having a
thiocarbonyl functional group was examined on the leaching of
synthetic chalcopyrite, covellite, bornite, and enargite.
Experiments were carried out in stirred reactors containing 40 mM
ferric sulfate solution at pH 1.8. 1 g of chalcopyrite or covellite
was added to the reactors along with an initial concentration of 2
mM of various thiocarbonyl reagents including Tu, TA, SDDC, ETC and
TSCA. The Cu extraction curves for chalcopyrite, covellite,
bornite, and enargite using all or a subset of the above reagents
are shown in FIGS. 18, 19, 20, and 21.
[0134] From FIGS. 18 to 21, it is clear that each of these further
reagents that have a thiocarbonyl functional group show a
beneficial effect in the ferric sulfate leaching of each of
chalcopyrite, covellite, bornite and enargite.
[0135] FIG. 22 summarizes the results of further stirred reactor
tests on chalcopyrite that additionally investigate urea and carbon
disulfide. These results confirm that, as expected, neither urea
nor carbon disulfide are effective reagents.
Example 5 FDS
[0136] The catalytic effect of leaching solutions prepared with FDS
on chalcopyrite, bornite, covellite, and chalcocite leaching was
determined in stirred reactor tests. All reactors contained 1.9 L
of ferric sulfate solution at pH 1.8 and total iron concentration
of 40 mM. 1 g of mineral samples was used in each reactor test. An
initial FDS concentration of 1 mM or an initial Tu concentration of
2 mM Tu was used.
[0137] The results from stirred reactor tests shown in FIG. 23,
panel a, and FIG. 23, panel b. demonstrate that FDS has comparable
efficiency to Tu in the leaching of each of chalcopyrite, bornite,
covellite, and chalcocite after 96 hours.
Example 6 Stepwise Closed Loop Bioleaching with Tu
[0138] A closed loop bioleach with Tu was conducted. 7 kg of ore
contain approximately 0.25% Cu content, mainly in the form of
CuFeS.sub.2 was leached at a flow rate of 1 L/day at an aeration
rate of approximately 300 mL/min.
[0139] The ore was pre-treated with sulfuric acid to leach oxides
(e.g. chalcanthite and basic copper salts) using sulfuric acid.
After the acid leaching period finished, residual solutions were
collected and replaced by a ferrous sulfate solution with nutrients
(40 mM FeSO.sub.4, 0.4 g/L magnesium sulfate heptahydrate and 0.04
g/L potassium dihydrogen phosphate, with pH adjusted to 1.6-1.8).
The ferrous and nutrients solution was flushed through the column
to establish a good habitat for bacterial growth. Inoculation of
bacteria showed an increase in the ORP from 274 mV to 550 mV within
48 hours. The solution used in this step and future steps was kept
circulating through the column, forming a self-sustaining
closed-loop system.
[0140] At this stage, the remaining copper source is mainly
CuFeS.sub.2. After the bacteria had survived in the column, Tu was
progressively added to the leaching solution. As discussed above Tu
is converted to FDS at a molar ratio of 2:1 in the presence of 40
mM Fe.sup.3+. Operating potential (ORP) was used as the indicator
for bacterial activity, and HPLC was used to monitor FDS content.
From day 0 to day 50, the leaching solution included 40 mM
Fe.sup.3+ with inoculated bacteria (with no Tu addition). From day
90 to day 98, a total of 1.878 g of Tu was progressively added,
upon which the HPLC analysis on the effluent showed that the FDS
was being maintained at approximately 1.5 mM, and no more Tu was
added.
[0141] As shown in FIG. 24, the ORP of the effluent was always
equal to or higher than the influent, indicating that bacteria were
actively oxidizing Fe.sup.2+ to Fe.sup.3+. The FDS contents were
analyzed by HPLC, showing that approximately 1.5 mM of FDS
(equivalent to 3 mM of Tu added) existing in the solution phase
without any precipitation being observed. Therefore, it appears
that 1.5 mM FDS (3 mM Tu equivalent) may be used in the solution
without precipitation of ferric.
[0142] The results of closed loop leaching test are shown in FIG.
25. From day 0 to day 50, bacteria were able to maintain high
activity and oxidize Fe.sup.2+ to Fe.sup.3+. However, with the
constant flow rate (1 L/day), the leaching rate was only 1.97 mg
Cu/day for the first 50 days. Addition of Tu starting on day 90
increased the Cu extraction rate to 6.54 mg/day, which remained
constant after day 98. This indicates that the reagent did not
undergo decomposition and remained effective in the closed-loop
system.
[0143] While specific embodiments of the invention have been
described and illustrated, such embodiments should be considered
illustrative of the invention only and not as limiting the
invention as construed in accordance with the accompanying
claims.
* * * * *